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By Edward Keller

GOLD MOVEMENT ON SURFACE OF AURIFEROUS COPPER WHEN LATTER IS SUBJECTED TO OXIDATION SOME years ago the writer's attention was called to the fact that rolling-mill scales from auriferous copper do not have the gold content proportional to the gold contained in the copper from which they are derived. I So, later, he made a few simple experiments, which are described herewith. In: order to test the gold movement on the changing surface of a copper plate under the influence of oxidation, plates 6 in. by 8 in. by I in. (15.2 cm. by 20.3 cm. by 2.5 cm.) of converter copper, refined, of 99.29 per cent. purity, and leaded copper with a content of 95.65 per cent. copper, were obtained. Ten cuts were made through the thickness of the plates along the longitudinal center and the silver and gold content ascertained in each of the 10 samples in both plates, as shown in the accompanying illustration. The assay results, recorded in Tables 1 and 2, show that the gold distribution is relatively uniform in both plates, while silver has greater variations, especially in the leaded plate. The plates subjected to oxidation were placed in a red-hot muffle for about 2 hr., two small scorifiers being used as supports. Arranged alongside each other, they nearly filled the width of the muffle, and so were practically under equal conditions of heat and draft. The accurate observance of the time was not possible, however, because of other work, but the time of the exposure of each plate in a pair was identical. At the end of each period, each plate was plunged into water and completely freed of scale, which was carefully collected, dried, weighed, and analyzed. Samples of the plates had also been taken of the molten metal and duly analyzed. Table 3 contains the analytical data for the two plates and their several sheets, or layers, of oxide scales. These figures show that in the oxidations of the converter copper the silver is fairly uniform, while

Jan 1, 1919

By C. H. Loux

This paper deals primarily with the advantages realized from the adoption of methods more suitable than those previously used. The improvements noted include: (1) modification of practice in shrinkage stoping, (2) use of sublevel caving over grizzlies, (3) improved haulage and loading, (4) use of contract wages wherever possible. The Washington mine is on Jonestown Hill, one mile south of Oxford, in northwestern New Jersey. It is producing 16,000 tons of concentrates a month, which are being shipped to the Alan Wood Steel Company's furnaces at Swedeland, Pennsylvania. History The first settlement in this region is reported to have been in 1735, and the magnetite ores were recognized immediately. The first furnace, built in 1741, began operation in 1743. During the Revolutionary War the output was chiefly used for making cannon balls, which, owing to the sympathies of the community, without doubt found their way into the hands of the British. This old furnace is still standing. Several important veins were soon developed, including the Martin, Strayley, McKinley, Adelaide, Welsh and Lanning. Mining was carried to a depth corresponding to the capacity of the pumps available. When these became inadequate, the workings were abandoned, and another vein was opened up. The Washington vein was probably discovered about 1850, and the first shaft started in 1854. In 1879, William H. Scranton made the first complete magnetic survey of the district and serious exploration of the deposit was begun. Rather indifferent success attended the earlier efforts of the many companies that attempted to mine and smelt the ore. In 1929 the property was leased by the Warren Foundry and Pipe Co. to the present operators, and the modification of practice then inaugurated is described in this paper. Geology These veins, like most eastern deposits of magnetite, are formed from, or are contemporaneous with, pegmatite intrusions into a gneissic country rock. The ore therefore follows the banded structure of the gneiss,

Jan 1, 1934

By Gerald Sherman

Mine fires are always dangerous and are frequently accompanied by loss of life during the period of confusion which is apt to follow their discovery. In metal mines, fires may result from the accidental ignition of combustible material, or they may be of spontaneous origin resulting from the rapid decomposition of sulphide. They occur frequently enough to demonstrate the need of some special equipment and a general plan for controlling them, in addition to measures for their prevention. Nearly all the large copper mines have had fires, and the methods of fighting them have been well described in recent papers, notably by C. L. Berrien,1 R. E. Tally,2 and others in discussion. The Copper Queen mine has many shafts, which are connected by underground workings. The deeper parts are ventilated by fans, but in the shallower portions natural ventilation through numerous connecting raises or shafts is sufficient. There have been several fires in the past few years, of which all are extinguished but one, which is under control. These notes cover particularly the preparations that are being made to check a new fire, partially or temporarily, and to facilitate the escape of workmen in the interval before a definite plan of fighting it can be established. A fire may occur in a shaft or station, or in some working farther away. In case of fire in an upcast shaft, the gas would escape to the surface and the immediate result would probably not be dangerous to workmen, as they could escape by other shafts. In a downcast shaft, however, the gas would be distributed very rapidly throughout a large portion of the mine, and men might easily be trapped and asphyxiated. A fire in mine workings, drift, raise, or stope, would be dangerous in a degree, but the draft there is likely to be less strong than in a shaft and probably a smaller area would be invaded by the fire gases. At the Copper Queen mine, in all drifts connecting with individual

Jan 1, 1918

By A. J. Miller, G. R. Brophy

THE Grossmann principle1 for the calculation of hardenability of steel from composition is attractive because of its simplicity. It postulates that the hardenability of a steel for any particular grain size may be expressed in terms of an ideal critical diameter, which is the product of an ideal critical diameter for a "pure" iron-carbon base multiplied by a succession of factors for each element of composition contained, including the incidentals. It infers that the multiplying factor for a given quantity of an element is constant for all combinations of composition. Grossmann has stated that in the great majority of cases the experimental values of hardenability are found to be well within 10 per cent of the calculated values. The validity of the scheme has been confirmed in principle by Crafts and Lamont2,3 and by Kramer et al .4 Their results differ from the original in the magnitude of the effect of some elements but, since they accepted the product concept and followed its direction, it is to be expected that similar results would be obtained. Crafts and Lamont, in their second paper, recognized that the factor relations for manganese and nickel were not continuously linear, but beyond certain critical amounts increase sharply. Austin, Van Note and Prater5 demonstrated nonlinear relations between hardenability and alloy contents. Their curves for chromium, silicon, manganese and nickel were all convex upward to show decreasing incremental effects. They concluded that Grossmann's factors underevaluate the hardening effect of low percentage additions of these elements, and overevaluate the larger additions. [ ] More recently W. Steven6 has found that the factor method fails when applied to complex steels containing two or three principal alloying elements other than carbon. Multiplying factors derived from his results bear a complex relation to composition and are not constant for a given amount of an element as the base composition varies. C. A. Liedholm7 found that, using either the Grossmann or the Crafts-Lamont multiplying factor values, not only were

Jan 1, 1945

By R. Janes, F. Everard

Pilot plant studies have been conducted on a variety of iron ores of differing composition and grain size to test their amenability to dry autogeneous grinding and dry magnetic concentration. A general description of the pilot plant and its equipment is given with particular reference to the Aerofall mill and the dry magnetic separators developed by the Ontario Research Foundation. Test results are reviewed in detail with emphasis on power consumption, mill speed, the use of ball charges, types of ores, moisture content, separator speeds, and design. In 1957 the Ontario Research Foundation installed a pilot plant (Fig. 1) for the study of dry autogenous grinding and dry magnetic concentration of iron ores. A 61/2-ft Aerofall mill was installed as the dry autogenous crushing and grinding medium. The Cavanagh fast drum and the Double Helix, two dry magnetic separators developed at Ontario Research Foundation, were also installed. Iron ore samples of 10 to 150 tons have been investigated. These ores varied widely in mineralogy and grain size. Taconite (-325 mesh), fine-grained magnetite (200 to 325 mesh), and medium-grained magnetites (65 to 150 mesh) have been treated with varying degrees of success. Lately, considerable research has also been conducted on the metamorphosed magnetite-hematite complexes of the Labrador Trough. Operational and metallurgical data have been compiled from these tests. The effects of ore type, air flow, mill speed, and ball charge on mill operation are discussed. Metallurgical data are also presented to illustrate the operation of the dry magnetic separators. PILOT PLANT DESCRIPTION Fig. 2 depicts the normal flowsheet used. Minus 6-in. ore from a vibrating feeder is fed to the 6%-ft Aerofall mill powered by a 50-hp 550-v induction motor. The main blower, a 25-hp 550-v fan rated at 7000 cfm, draws air through the mill. Finely ground ore is conveyed by the air stream to a horizontal classifier, where expansion and a change in direction cause the +I50 mesh fraction to fall onto a vibrating

Jan 1, 1962

By H. E. Nold

THE raw materials of the ceramic industry are mostly clays. This raw material is ground, water is added and the mixture pugged into a moist, plastic, rather stiff mass. From this mass the desired units of manufacture are made (bricks, hollow building tile, sewer pipe, etc.). These units are dried; then placed in kilns and heated until the desired degree of vitrification is obtained.1 For firing ceramic ware, two distinct types of kilns are used; namely, periodic and continuous tunnel. PERIODIC KILNS Periodic kilns are those ordinarily seen in a brick plant. They are either round or rectangular. The raw, dried, formed ware is piled or set in the kilns to depths frequently exceeding 10 ft. The kiln is heated until the ware is matured, then gradually cooled sufficiently to allow men to remove the ware; then the cycle is repeated. The time duration of the cycle may be more than two weeks. In order that heat can be distributed uniformly throughout the pile or setting, the units are placed so as to leave open spaces or flues through-out the pile for the passage of the hot gases. Each kiln is provided with a number of fireboxes (usually 10 or 12) spaced equally in the outside wall. The fireboxes may be equipped with flat or inclined grate bars or they may have no grates at all (dead bottom). Down-draft kilns are most used. In these the hot gaseous products of combustion from the fireboxes are deflected upward by bag walls just inside the fire zone. From the top of the kiln the gases are drawn down through the setting, off through holes in the kiln bottom, and conducted to a stack or stacks. In up-draft kilns the fireboxes are usually set somewhat lower. The hot gases are drawn inward and upward through the ware and pass off from the top. Up-draft kilns find their greatest use in the firing of common brick. In either case the problem is to maintain, as nearly as possible,

Jan 1, 1933

By A. Ahmadieh, J. E. Dorn, Jack Mitchell

A detailed investigation of the disloccrtion mechanisms controlling prismatic, slip in a solid solutions of magnesium containing up to 15.9 at. pel Li revealed that low-temperature prismatic slip is controlled by the Peicrl's mechanism for all alloy compositions. At higher temperatures slip in the 7.9 at. pet Li alloy was found to follow the diclates of the cross-slip mechanism from basal to prism planes, while the higlier-lilhium alloys exhibited an athermal dejormation process which might be due COARSE-GRAINED polycrystalline aggregates of distilled magnesium exhibit very limited ductility at low temperatures.' The ductility of polycrystalline magnesium, however, is much improved by solid-solution additions in excess of about 8 at. pct to shot-range ordering. The major effect of lithium additions on the low-temperature Shear] stress. for prismatic slip appears to be due to its efftct in lowering the Peierl's stress whereas (1 increase in the shear stress with increased lithium conlent at higher temperatures might he due to short-range order strengthening. Additions rip to 7.9 (at. pel Li appear to have only minor effects on the slackiug-jaidt energy. Li.2 Although the relative amounts of basal slip and twinning are not materially altered by such alloying, a substantial change takes place in the relative amounts of prismatic slip. Whereas prismatic slip is negligible and limited to regions of high stress concentrations in the vicinity of the grain boundaries in pure magnesium, extensive prismatic slip over entire grains occurs in the higher lithium content alloys. Furthermore, as the lithium content is increased above 8 at. pct, the flow stress for polycrystalline aggregates decreases. Comparison of the single-crystal data on pure magnesium for basal slip by Sheely and Nash3 and those for prismatic slip by Flynn, Mote, and Dorn4 indicates that the critical resolved shear stress for prismatic slip is much above that for basal slip.

Jan 1, 1965

By R. E. Zimmerman

WITH coal-preparation plants becoming more complex and containing widely diversified types of equipment and processes, it is necessary that considerable control be exercised in their operation. Rising mine costs make it imperative that close check be made of plant efficiencies to assure a minimum loss of coal to the refuse bank. Further, much of the coal sold today is sold on a competitive specification basis, and this makes it necessary to provide frequent analyses and inspections. Many ' plants are required to operate on a two or three-shift basis, and as a con- sequence an efficient organization is required to eliminate down time and provide for smooth low-cost operation. ORGANIZATION Probably the most important item for successful operation of a preparation plant is a capable, properly organized staff, with the individual workers properly informed and well trained for their particular jobs, clear-cut lines of authority among the supervisors and in general a standard operating procedure for most conceivable conditions that may occur. The idea of placing the older, partially disabled and generally incompetent coal miner in the cleaning plant, as was the practice on the earlier tipples, is entirely inadequate in a modern complex cleaning plant. The new plants require intelligent, alert, and reasonably educated men. With perhaps millions of dollars invested in an installation, it is poor business to use other than a well organized, well trained, and properly supervised force. Organizations naturally must vary with the size of the plant and local conditions; whether it is a central cleaning plant at a distance from the mine organization or one of several cleaning plants operated by the company. Smaller operations can be operated by a washery foreman, under the supervision of the mine superintendent or perhaps the "outside" superintendent. Large installations should have a plant superintendent either responsible to the mine superintendent or preferably, where possible, responsible to a preparation manager or engineer. A typical organization chart for companies operating in the anthracite region of Pennsylvania is shown in Fig. 1. Breaker (preparation plant)

Jan 1, 1950

By Richard A. Arterburn

For many years hydrocyclones, commonly referred to as cyclones, have been extensively utilized in the classification of particles in comminution circuits. The practical range of classification for cyclones is 40 microns to 400 microns, with some remote applications as fine as 5 microns or as coarse as 1000 microns. Cyclones are used in both primary and secondary grinding circuits as well as regrind circuits. The information given in this chapter is intended to provide a method, at least for estimating purposes, of selecting the proper number and size of cyclones and to determine the proper level of operating variables. Generally, it is recommended that cyclone suppliers be consulted for sizing confirmation. Sow cyclone suppliers employ digital computers to aid in the sizing and selection of cyclones. CYCLONE DESCRIPTION AND BASIC OPERATION Figure 1 shows a cutaway view of a typical cyclone. During operation, the feed slurry enters the cyclone under pressure through the feed pipe into the top of the cylindrical feed chamber. This tangential entrance is accomplished by two types of design, as shown in Figure 2. Since the majority of research has been done with the involuted type, the graphs and relationships shown may not be strictly applicable to other designs. As the feed enters the chamber, a rotation of the slurry inside of the cyclone begins, causing centrifugal forces to accelerate the movement of the particles towards the outer wall. The particles migrate downward in a spiral pattern through the cylindrical section and into the conical section. At this point the smaller mass particles migrate toward the center and spiral upward and out through the vortex finder, discharging through the overflow pipe. This product, which contains the finer particles and the majority of the water, is termed the overflow and should be discharged at or near atmospheric pressure. The higher mss particles remain in a downward spiral path along the walls of the conical section and gradually exit through the apex orifice. This product is termed the underflow and also should be discharged at or near atmospheric pressure.

Jan 1, 1982

By Ernest Gayford

When the United States Government started buying gold at varying prices per ounce, set by the President and the Treasury Department, the gold miner found himself facing new conditions, as gold became a commodity with a fluctuating market price. While it was self-evident that the miner was to receive more for his gold than the old established price of $20.67 per ounce, there was an element of uncertainty, which made it difficult, and perhaps dangerous, to decide to invest his time and money to produce gold, especially if the profit would represent only the difference between $20.67 an ounce and the new price as set from day to day by the Government. When, however, on Jan. 30, 1934, the United States went off the gold standard, the price of newly mined gold being rather definitely set at $35 per ounce, with some assurance that the price would be maintained at or near this figure for a long time to come, the gold miner was justified in revaluing his property on the basis of the new price, and the prospector was offered inducement to spend his time and energies hunting for new mines, or going back to old locations previously abandoned as of doubtful value. It is interesting to see what this new price means to the producer of newly mined gold, and how much of the increase is passed on to the producer. It will be found that, as with the base metals, the nearer he can get to producing a finished product himself the greater his reward will be; the net returns for bullion being necessarily higher than for the same amount of gold shipped in concentrates to the smelter. Bullion can be shipped to the United States mints or Government assay offices under the following terms and conditions. An affidavit must accompany the shipment, stating, among other things: 1. The source of the gold; whether ore, tailings or placer. 2. The tons of ore, tailings or cubic yards of gravel from which the shipment was recovered.

Jan 1, 1935

By G. W. Jarman

THE purpose of this paper is to present briefly a description of some of the special methods of separation or concentration, either singly or in combination with others, and to give certain operating data and examples. Explanation of the usual methods of large-production mill treatment, such as flotation or wet tabling, or of chemically treating industrial minerals, is not included in the scope of this paper, although one of these methods might be, in certain instances, augmented by one or more of the special methods that will be enumerated herein. Description of these special methods will be centered on magnetic separators, electro- static separators and air flotation or air tables. MAGNETIC SEPERATORS In general, separating is achieved by exerting upon a particle some kind of force opposed to gravity, either directly or at an angle. For example, in the Wetherill or cross-belt type of magnetic separator some particles are lifted vertically from the others because they have been exposed to a force to which they respond more readily than particles of lesser magnetic susceptibility. Of course, there is a limitation to this vertical rise, for the particle must be of a high enough susceptibility to receive a magnetic force greater than the 980 dynes per gram of gravitational urge. To overcome this limitation, the induction type of magnetic separator was developed, so that it is only necessary to instill, say, 30 to 60 dynes per gram instead of the 980, to cause a deflection away from the vertical in falling, and that deflection can be used advantageously when falling by inserting a dividing knife-edge. The induction magnetic separator (Fig. 1) has become a fairly well known method of concentrating and purifying. For example, abrasive grain was among the materials that did not respond to the previously best known type of magnetic separator, such as the Wetherill type, and the development of the induction magnetic separator in the plant of a

Jan 1, 1938

By H. H. Uhlig

IN our study of the mechanism of and the prevention of corrosion in stainless steels, it was considered of fundamental importance to obtain knowledge of the surface structure of such alloys. If corrosion resistance is due to a self-healing oxide layer, which like a paint film protects the underlying metal, it would be obviously relevant were the effort made to make the film less permeable and more adherent; for example, the surface might be expected to respond to an anodizing process, which is found successful in increasing the film thickness and corrosion resistance of aluminum. However, if passivity and corrosion resistance of stainless steels are due to another cause, the attack of the problem would follow another path. One of the most effective techniques for studying passivity and corrosion tendency is the measurement of the electrochemical potential. Not only can one determine by potential measurements whether a metal is active or passive, but also one can determine the degree of passivity and rate of transformation of passivity with time. Chemical media that build up or destroy passivity when in contact with the steel can also be classified by potential measurements. Something can be learned, too, of the effect of alloy additions such as molybdenum to 18-8, known to decrease corrosion of the alloy in chloride solutions. A study of alloys in previous investigations has sometimes included data on the electrochemical potentials. Tammann and Sotter1 measured the electrochemical potentials of alloys of iron-molybdenum and iron-chromium and reported the critical concentration of molybdenum or chromium at which passivity initiates. Strauss2 made similar measure-ments for the iron-chromium system. The electrochemical potential' has also been used to indicate carbide precipitated 18-8. In 18-8, heat-treated to precipitate carbide, active metal exists at the region of the grain boundaries, contrasted with passive metal within the grains, the active metal accounting for a resultant active potential. Several

Jan 1, 1939

By A. S. Adams

A GLANCE at the list of papers1 that have been published since 1920 on the general subject of flotation suggests the variety of ideas that exist regarding the underlying cause of the phenomenon.' Among these ideas, we find several coherent theories, but none that adequately and completely explain the phenomenon. Ralston2 and Bains3 suggested an electrical theory which has to do with the attachment of mineral particles to a bubble by the attraction of unlike electrical charges. This theory has its root in the fact that a bubble of air in pure water is electrified in such a way that the surface is negatively charged while the body of the water is positively charged. This result, which was first discussed by Helmholtz, also obtains when the water contains anything except a high concentration of hydrogen ions, or ions which are trivalent or tetravalent. The theory also holds that the mineral particles have positive charges on their surfaces, and so there is attachment between bubble surface and mineral particles by the old law of electrostatics, which states that unlike charges attract. CONTACT ANGLE AND OIL ATTRACTION The contact angle-the angle formed by a fluid-fluid boundary when both are in contact with a solid-has been indicted several times for complicity in the phenomenon of flotation. This angle is supposed to indirectly measure the surface energy of a solid. Thus if we have two fluids in contact with a solid, we may express the difference between the two fluid-solid tensions as equal to the fluid-fluid tension times the cosine of the angle between the two fluids.4 If the two fluids are in contact with the solid, this is, no doubt, an excellent way of establishing relative surface energies. However, in flotation, we deal with substances which are not wetted by water. What assurance have we, then, that the angle

Jan 1, 1928

By Edward Keller

(Canal Zone Meeting, November, 1910.) UNDER the title, Labor-Saving Appliances in the Works-Laboratory, I published a paper1 in which was described how multi-manipulations, in a works-laboratory, and in the furnace-room of an assay-laboratory, can be condensed into single manipulations by applying the proper mechanical devices. For example, I now deposit in, or withdraw from, a muffle a set of 48 cupels as one unit. The front row. (8) of these. are blanks or heaters; the other 40, when they have attained the proper temperature in the muffle, are charged simultaneously with the 40 lead-buttons. Three manipulations here accomplish an operation which by the generally customary method requires 136, and this number is only limited by the size of muffle and cupels. This operation is an enlargement of what I have already described in my former paper, and Fig. I shows the improved implements. In the paper referred to I described a gold-silver bead parting-bath, which is the last of the devices used in regular sequence in practice. This bath was designated as being convenient, but was not a labor-saver properly. Fig. 2 of the present paper shows a new device. Instead of the original tray, there are now sectional holders for the test-tubes, each having a wooden handle on either end, so that the holders may be removed from the boiling bath, and the acid or water poured off from each set of tubes (in this case, 7) without waiting to cool. The tubes, held in place by clutches, as shown in Fig. 3, rest in holes in the base-strip, having a smaller diameter than the tubes. Each holder is stamped at either end with a number, so that the bath becomes further useful by permitting several men to use it at the same time without interference. Apart from the handles, the holders are made of sheet-copper. ' Trans., xxxvi., 3 to 18 (1906).

Aug 1, 1910

By L. A. Delano

DURING 1916, the St. Joseph Lead Co. milled 2,505,670 tons of ore. This is a daily operating average of 7855 tons. The economic concentration of such a large tonnage necessarily requires a plant equipped with modern machinery and operated efficiently. The milling process used in producing this output has been a gradual development through the long period of years that the company has been in existence. HISTORICAL SUMMARY OF MILLING PRACTICE Ore was first mined at Bonne Terre, Mo., in 1864. The early milling practice was crude. It consisted of breaking the ore into small pieces by hand, pulverizing in a Blake crusher and a pair of Cornish rolls to approximately minus 6 mm., and jigging in the old-fashioned hand jigs. Only the coarse heavy galena was saved, the gangue being shoveled off and discarded. The output for the first year under this method was only 240 tons of pig lead. The first mill of the company was a combination of extensions and additions built from time to time as the mines developed. It was a museum of early mulling machinery, such as jigs, percussion tables, huddles, dolly tubs, log washers, etc., following no definite flow sheet. The chats, or tailings, were hauled to waste dumps by mule and ox teams, as much as 500 tons per day being moved in this manner, until a railroad was built in 1880. This mill was burned in 1883, at which time the annual output had reached approximately 7300 tons of pig lead. A new mill of 500 tons capacity was designed and built by C. B. Parsons. This mill has been admirably described by H. S. Monroe.1 The ore was crushed in Blake crushers, Cornish rolls, through 6-mm. screens, and treated by jigs of the Parsons type, percussion tables and log washers, or trunking machines, for cleaning the galena. The middlings were recrushed in rolls and treated on Harz jigs. The mill was operated without change, except later installing Gates crushers and en-

Jan 9, 1917

By M. Van Siclen

THE outstanding fact in connection with dust disease in the United States at present is the growing recognition of its seriousness by state officials and by the more progressive operators of mining, metallurgical and allied industrial plants. The same may perhaps be said of labor in some of its better organized centers, but the average independent non-union workman continues, .as of yore, to suffer the toll of disability and even of death from dust inhalation without realization of the source of his ill-being. Progress in the betterment of the handling of sufferers from dust disease and of the conditions creating dust hazards in mining and allied industries is pursuing a course similar to that of the now widespread campaign against accidents, in which the risk is more obvious, more severe, and was naturally recognized sooner. The coming on of dust disease is generally relatively slow and its immediate effects not infre-quently masked by or attributed to other diseases. Marked exceptions to this general rule have occurred in districts where the dust is particularly pernicious, such as the Rand and Joplin; yet such exceptions, unhappy in themselves, opened the eyes of medical men, and of those in other localities charged with the responsibility for the welfare of industry, to the much more widespread if less spectacular hazard existing from the inhalation of dust in all rock-excavating operations and all industrial operations creating inorganic dust, whether it be crushing and grinding, sand-blast-ing, granite-cutting, or what not. Advancement to date in the United States in prevention and treat-ment has been slow, yet the record shows that quantitative methods for determining the degree of hazard and means for decreasing it were being developed as far back as 25 years ago; and that more rapid progress may be looked for from now on. In 1908, in this country; F. L Hoffman?(1) discussed the mortality from consumption in dusty trades, especially brass foundries; and in 1913 E. R. Hayhurst(2) described the effects of inhalation

Jan 1, 1933

By P. BOURY

THIS account of a contrivance which has been found serviceable in practice may be of interest to engineers, and' especially to those engaged in hydraulic mining. In that process, one feature which seems at first sight to possess little significance, is highly important in its effects, namely, the regular supply of water to the " giants." The chief condition of this regularity is an even flow in the pipe-line from the penstock to the nozzle. This, of course, requires that the penstock be kept full. Any considerable surplus of water, overflowing at the penstock, is a simple waste, unless it be otherwise utilized in the sluices. At all events, it does not affect the delivery of the nozzles. But a deficiency at the penstock may admit air into the pipe; and this may produce hydraulic recoil, injuring the line by incessant vibrations, giving rise to leaks, and even to blow-outs, and (when a partial vacuum exists at the time of the fracture the collapse of the pipe. An accident of this sort in our main pipe-line, occasioned partly by a settling of the ground and partly by the presence of air in the pipe, led to the employment of the device here described. The line referred to is made of 30-in. pipe, successively reduced to 18 and 15 in. in diameter, for the supply of two giants. By reason of the causes above mentioned, the 30-in. pipe burst more than 1,400 ft. from the penstock, and, despite the air-valve, the 30-in. pipe was collapsed to such a degree that one length of it, next to the penstock, was squeezed together and positively swallowed by the main pipe, through which it traveled for 750 ft., piercing through a short elbow on the way. This remarkable result was undoubtedly the effect of the vacuum suddenly produced when the pipe burst.

Sep 1, 1906

By J. L. Bray

Lead has often been suggested as a protective coating for iron and steel. Such a protective coating should possess: (1) good adhesion, (2) durability, (3) ease of application, (4) freedom from pinholes, (5) good appearance and (6) low cost, but, unfortunately, lead coatings so far produced do not have all of these characteristics. Because of the necessity of carrying out the coating process cheaply, and because lead solutions have poor plating characteristics, most of the efforts have been directed toward the hot-dipping process rather than electroplating. The hot-dipping processes using lead have not been successful chiefly because of the lack of adherence between iron and lead. In any coating process involving iron a bond can be obtained only by the formation of an intermetallic compound, the formation of a solid solution, or mechanical means. Lead forms with iron two immiscible liquid solutions, and in the solid state no intermetallic compounds or solutions. It is necessary, therefore, to employ some other element or elements as a binding agent. Tin may be used as in terneplate, or copper in the form of a thin film of electrolytically deposited metal to which the lead will adhere. Because of its physical and chemical properties, lead as a coakkg medium would undoubtedly be limited in its use to sheet and wire and then only when the article is not subject to abrasion or wear. A very large tonnage of these products is being used under such conditions and a successful lead-coating process should be a valuable contribution to the science. Two other difficulties stand in the way of successful lead coating—that of poor appearance of the weathered surface and the tendency of the lead to form pinholes during solidification. Since lead is commonly accepted as cathodic to iron, these exposed areas are ones of accelerated corrosion. Use of Zinc.—The method under consideration involves zinc as a binding agent on both the steel sheet and in the molten bath. This

Jan 1, 1937

By Mark A. Ammon

The two most widely used methods of measuring hardness are the Brinell and the scleroscope. In the Brinell method a hardened steel ball is pressed into the steel under a definite load and the area of the resulting depression is measured. The load in kg. per sq. mm. of the area of the depression is taken as a measure of the hardness. Within certain limits the hardness number so determined is independent of the load or the size of the ball. In the scleroscope method a small weight carrying a diamond point is dropped through a glass tube upon the steel and the amount of rebound is taken as proportional to the hard-' ness. In both methods the scale of hardness is an arbitrary one; however, they agree in a comparative may, that is, a steel showing 50 per cent. increase in hardness by the Brinell test mill show approximately 50 per cent. increase in hardness by the scleroscope test. Yet upon consideration of the two methods it appears that, upon a steel which has been carbonized, this should not be the case ; for the scleroscope method gives almost exclusively the surface-hardness, and the condition of the material below the surface has but a slight bearing upon the hardness number. The results given by the Brinell method should, however, be the exact opposite; for the depth of penetration 'of a ball pressed into the steel would be dependent to a large extent upon the condition of the material below the surface. This is particularly obvious when it is considered that the size of the ball ordinarily used in the Brinell method is 1 cm. (about 8 in.) in diameter under a load of 3,000 kg., while the diamond point of the scleroscope is in the neighborhood of 0.01 in. across, fitted in a hammer of about 0.5 oz. weight and dropping from a height of 10 inches.

Jan 1, 1913

By Research Staff ? Verde Copper Mines

FLOTATION is now commonly practiced in alkaline ore pulps, yet little is known regarding the action of the alkaline solutions on the ore particles beyond the fact that films of oxidized material form. We wish to chronicle our discovery of the presence of soluble reducing agents in alkaline ore pulps of the United Verde, and present some data on lime consumption. During the grinding and flotation of a pyritic ore containing small amounts of ferrous sulfate and other soluble salts, an excess of lime was added to precipitate the salts, and filtered samples of the solution were titrated with standard permanganate solution. Even in pulps of high causticity an appreciable permanganate titration was always obtained, in spite of the fact that the solutions were shown to contain little iron. Investigation of the cause brought about the discovery of several previously unsuspected compounds in flotation pulps made from pyritic ore. Not only the reducing agents present had to be identified, but also the amount of dissolved oxygen, and a survey was finally made of the free oxygen content and reducing agents of the mill pulp in the concentrating mill at Clarkdale. A great many compounds might form, caused by oxidation of pyrite or other metal sulfides or by reaction between a lime solution and pyrite. The most probable compounds considered were calcium sulfides; polysul-fides, thiosulfates, and polythionates. Later it was discovered in the literature that in alkaline pulp polythionates do not exist and if introduced. are decomposed into sulfites and thiosulfates.1 Any precipitated sulfur is changed by alkali into thiosulfate and polysulfide. Sulfites and sulfides react to form thiosulfate, so that in the end it is possible to have in an alkaline solution only thiosulfate and an excess either of sulfide or sulfite. Confirmatory tests were made and it was finally established that the alkaline mill pulps made from the United Verde ore contained

Jan 1, 1929